Method for recovering platinum group elements

ABSTRACT

A method for recovering platinum group elements comprises charging an oxide raw material entraining platinum group elements, copper oxide, solid carbonaceous reducing agent and flux into a closed electric furnace equipped with at least one electrode for heating material and a roof for substantially shielding material charged into the furnace from the exterior atmosphere. The charged material is melted and reduced to form a layer of molten metal under a molten slag layer and enrich platinum group elements in this molten metal. The oxide raw material, copper oxide, solid carbonaceous reducing agent and flux are prepared as powdery or particulate materials, premixed and then charged into the electric furnace. The heat-melted material in the furnace is allowed to stand at a temperature of 1200° to 1500° C. for at least 5 hours, with the molten metal containing platinum group elements then being discharged outside the furnace.

TECHNICAL FIELD

This invention relates to a method for recovering platinum groupelements from a substance containing at least one platinum groupelement, such as spent petrochemical type catalyst, spent vehicleexhaust gas purification catalyst, and spent electronic circuit boardsor lead frames

BACKGROUND ART

Conventional methods for recovering platinum group elements from spentvehicle exhaust gas purification catalysts and other platinum groupelement-containing substances include, for example, the method ofextracting platinum group elements by using a solution of an oxidizingagent added acid such as aqua regia and the method of dissolving acarrier in sulfuric acid or the like and separating the undissolvedplatinum group elements. However, such wet methods have been found to beimpractical owing to inferior platinum group element extraction rate andto recovery and cost problems owing to the need to use large amounts ofacid for dissolving the carrier.

To replace the wet method, JPA-4-317423 teaches a dry method forrecovering platinum group elements. This dry method is a revolutionarymethod that recovers platinum group elements at high recovery rate andlow cost by absorbing platinum group elements contained in spent vehicleexhaust gas purification catalyst or the like into molten copper andconcentrating the result. JPA-2000-248322 teaches a dry platinum groupelement recovery method that is an improvement of JPA-4-317423.

The methods of JPA-4-317423 and JPA-2000-248322 are superior to the wetmethod in the point of enabling recovery of platinum group elements(hereinafter sometimes abbreviated as “PGM”) at high recovery rate andlow cost in a short period of time and have a particularly advantageousfeature in that the recovery rate of Rh in PGM is higher than by the wetmethod.

However, when a dry method is conducted in an electric furnace, i.e.,when a PGM-containing substance (spent vehicle exhaust gas purificationcatalyst or the like) and copper or copper oxide is melted by heating inan electric furnace in the presence of a flux component and a reducingagent and the PGM of the substance is absorbed into the molten metalliccopper, the following problem emerges. Specifically, during theabsorption of the PGM into the molten metallic copper, some amount ofPGM is also entrained by the electric furnace slag layer present as theupper layer, so that when the electric furnace slag is discharged, PGMalso flows out of the system as a result. This lowers the PGM recoveryrate.

On the other hand, when a dry method is conducted in an electricfurnace, the furnace rises to a high temperature that makes it necessaryto protect the furnace body with heat-shielding means or cooling means.Furnace body protection methods of this type include lining the innerwall of the furnace with a refractory material and cooling the furnacebody by blowing air onto the outer wall of the furnace, and installing awater-cooling jacket or an oil-cooling jacket. However, the inner wallrefractory at the height level where the molten slag is present and/orat higher levels than this sustains severe damage, and when the damagebecomes excessive, the operation has to be discontinued to carry outrepair work. In a refining furnace that uses a carbonaceous reducingagent to obtain a metal melt by reducing an oxide raw material chargedinto the furnace, the refractory inside the furnace is particularlysusceptible to damage and restoring it requires a large amount ofexpenditure.

DISCLOSURE OF THE INVENTION

The present invention was therefore accomplished with the object ofovercoming the aforesaid problems by providing a method and equipmentfor recovering PGM by the dry method from PGM-containing substances withgood operability at high yield.

As a method for recovering platinum group elements for achieving thisobject the present invention provides a method for recovering platinumgroup elements comprising:

charging an oxide raw material entraining platinum group elements,copper oxide, solid carbonaceous reducing agent and flux into a closedelectric furnace equipped with electrodes for passing electric currentthrough and heating material charged in the furnace and a roof forsubstantially shielding material charged in the furnace from theexterior atmosphere,

melting and reducing the charged material by the electric currentheating with the electrode, thereby forming a layer of molten metalunder a molten slag layer and enriching platinum group elements in thismolten metal,

which method is characterized in that:

the oxide raw material, copper oxide, solid carbonaceous reducing agentand flux are all prepared as powdery or particulate materials;

these powdery or particulate materials are premixed and then chargedinto the electric furnace; and

the molten metal enriched in platinum group elements is dischargedoutside the furnace after undergoing a standing step in which theheat-melted material in the furnace is allowed to stand at a temperatureof 1200-1500° C. for at least 5 hours.

Here the slag oxides formed in the furnace are preferably controlled tothe content ranges of: Al₂O₃: 20-40 wt %, SiO₂: 25-40 wt %, CaO: 20-35wt %, and FeO: 0-35 wt %. Moreover, at this time, the content ranges ofthe slag oxides are preferably controlled by the steps of: analyzing andascertaining beforehand the amount of at least one among the oxides,preferably all of the oxides, of Al, Si and Fe contained in the oxideraw material entraining platinum group elements, and regulating thecomposition of the flux components before charging them into the furnacein accordance with the contents of these oxides. The molten metalobtained by implementing this method can, after being discharged fromthe furnace, be transferred to and oxidized in another furnace andseparated by difference in specific gravity into an oxide layer composedmainly of copper oxide and molten metal composed mainly of metalliccopper in which platinum group elements are enriched. The separatedoxide layer composed mainly of copper oxide can be reused as theaforesaid copper oxide raw material charged into the electric furnace.

In accordance with the present invention there is further provided amethod for recovering platinum group elements comprising:

charging an oxide raw material entraining platinum group elements,copper oxide, solid carbonaceous reducing agent and flux into a closedelectric furnace equipped with electrodes for passing electric currentthrough and heating material charged into the furnace and a roof forsubstantially shielding material charged into the furnace from theexterior atmosphere,

melting and reducing the charged material by the electric currentheating with the electrode, thereby forming a layer of molten metalunder a molten slag layer and enriching platinum group elements in thismolten metal,

which method is characterized in that:

as the electric furnace there is used an electric furnace whose wallenclosing up to a height level of the molten slag layer generated in thefurnace is constituted of an iron shell and a sheet of flowing water isformed on the outside of the iron shell to descend in contact with itsouter surface.

The sheet of flowing water descending in contact with the outer surfaceof the iron shell can be formed by evenly distributing water of aprescribed water head pressure from a header installed outward of anupper part of the furnace wall toward the whole outer periphery of theiron shell at the same height level.

BRIEF DESCRIPTION OF THE DRAWINGS

FIG. 1 is schematic sectional view of an essential portion of anelectric furnace in accordance with the present invention.

FIG. 2 is a schematic perspective view of a portion of the wall of anelectric furnace in accordance with the present invention.

FIG. 3 is a system diagram of electric furnace cooling water processingsystem in accordance with the present invention.

PREFERRED EMBODIMENTS OF THE INVENTION

FIG. 1 shows an example of an electric furnace suitable for implementingthe present invention. The electric furnace of FIG. 1 is a closedelectric furnace equipped with electrodes 2 for passing electric currentthrough and heating charged material 1 in the furnace, and a roof 3provided on a furnace body 4 for substantially shielding the chargedmaterial 1 from the exterior atmosphere. The charged material 1 iscomposed of a mixture of oxide raw material, carbonaceous reducing agentand flux. The charged material 1 is melted by electric current passingthrough it from the electrodes 2 to form a molten slag layer 5, whilemolten metal 6 generated by reducing the oxides simultaneously poolsbelow the molten slag layer 5. The generated molten slag layer 5 andmolten metal 6 are suitably discharged outside the furnace through aslag tapping hole 7 and metal tapping hole 8. Although not illustrated,gas generated inside the furnace is discharged to outside the furnacethrough an exhaust gas passage. Symbol 9 in FIG. 1 indicates a materialloading chute.

In the present invention, the wall of the electric furnace enclosing upto the height level of the molten slag layer generated in the furnace isconstituted of a iron shell 10 and a sheet of flowing water 11 is formedon the outside of the iron shell 10 to descend in contact with its outersurface. In the illustrated example, the wall of the furnace body 4 (thepart of the furnace other than the bottom and the roof 3) is constitutedthroughout by the iron shell 10. In other words, the furnace wall isformed by the iron shell 10 not only up to the height level where themolten slag layer 5 is present but also from thereon up to the regionwhere the wall contacts the roof 3. FIG. 2 illustrates this statediagrammatically.

In FIG. 2, the entire furnace wall is formed of the cylindrical ironshell 10. When the furnace wall is formed of the iron shell 10 in thismanner, an annular header 12 is installed outward of an upper part ofthe iron shell 10 to encircle the iron shell 10 at a certain spacingtherefrom. Similarly, an annular trough 13 is installed on the outsideof the lower part of the iron shell 10 to encircle and make contact withthe iron shell 10. Water is evenly sprayed from the annular header 12toward the whole outer peripheral surface of the iron shell 10, therebyforming a curtain like sheet of flowing water that descends along theouter surface of the iron shell 10 to be received in the annular trough13.

This continuous sheet of flowing water 11 (FIG. 1) formed on the outersurface of the iron shell 10 continuously cools the inner surface of theiron shell 10 so that molten slag generated inside the furnace forms aself-coating 14 of a regular thickness. The self-coating 14 not only isformed at the level where the molten slag layer 5 is present inside thefurnace but may also be similarly formed on the inner surfacethereabove. Therefore, even if no furnace wall lining is installed atthe time of furnace construction, the function of a lining is producedup to the level of the molten slag layer 5 chiefly by the formation ofthe self-coating 14 composed of solidified slag formed on the innersurface of the iron shell 10. If a portion of the self-coating 14 shouldbe damaged in the course of operation, the damaged portion isautomatically restored in the course of operation, so that, differentlyfrom in the case of a refractory lining, a need to carry out repair workon damaged portions seldom arises.

FIG. 3 diagrammatically shows the cooling water system for forming thecontinuous sheet of flowing water 11 on the outer surface of the ironshell 10. As shown in FIG. 3, the header 12 is installed around the ironshell 10 with its axis oriented horizontally and its side facing theiron shell 10 is formed with a slit-like nozzle opening 15. The annularheader 12 is supplied with cooling water of a prescribed water headpressure so that cooling water flows out from the slit-like nozzleopening 15 horizontally in a sheet-like manner toward the iron shell 10to form a steady annular flow that falls along a parabolic trajectory.The distance between the iron shell 10 and the slit-like nozzle opening15, the size of the nozzle opening, the pressure in the header, thequantity of supplied water and the like are determined so that thissteady annular flow makes contact with the cylindrical surface of theiron shell 10, namely, so that the annular flow spreads out naturally tocover the shell 10 completely, without splashing off the surface of theiron shell 10 and without separating from the iron shell 10 during itsdescent after reaching the iron shell 10 or forming branching orconverging flows. For this purpose, the temperature, quantity andpressure of the cooling water are controlled as explained in thefollowing.

The sheet of flowing water 11 that flows downward along the outersurface of the iron shell 10 to be received in the annular trough 13 isfirst pooled in a storage reservoir 16. The reservoir 16 is suppliedwith makeup water from a makeup water conduit 17 in an amount equal tothe evaporation loss. The hot water in the reservoir 16 is pumped to asprinkler 20 of a cooling tower 19 by a pump 18. The hot water is cooledby heat exchange with external air at the cooling tower 19. The cooledwater is pumped up to a water head tank 22 by a metering pump 21 andthen supplied into the annular header 12. At this time, the amount ofwater sent from the water head tank 22 into the annular header 12 iscontrolled so as to keep the water level in the annular header 12 level.Since constant pressure cooling water discharges from the slit-likenozzle opening 15, the continuous sheet of flowing water 11 is steadilyformed on the surface of the iron shell 10. Owing to the steadyformation of the sheet of flowing water 11, the self-coating 14 ismaintained on the inside of the iron shell 10 at a constant thickness.

The method for recovering platinum group elements from used anddiscarded catalyst and the like will now be explained. The electricfurnace explained in the foregoing can be used to implement the recoverymethod.

The “platinum group element-containing substance to be processed” astermed with respect to the present invention is, for example, typicallya spent and discarded petrochemical type catalyst containing platinum,palladium and the like or a spent and discarded vehicle exhaust gaspurification catalyst containing platinum, palladium and optionallyrhodium and the like, but also includes rejected products, scrap and thelike occurring in the process of manufacturing such catalysts, as wellas used electronic circuit boards, dental products, lead frames and thelike that contain palladium and the like. Such substances for processingthat contain at least one platinum group element are ordinarily in aform of having a small amount of platinum group elements carried on ametal oxide or ceramic.

The present invention is basically constituted of the steps of charginginto a refining furnace and melting so-defined substances for processingthat contain at least one platinum group element, together with a coppersource material containing copper oxide, a flux and a carbonaceousreducing agent, sinking a molten metal layer of primarily metalliccopper below a formed molten slag layer of primarily oxides, andenriching the platinum group elements in the molten metal sunk below. Atthis time, the present invention adopts some distinctive features asfollows:

-   1. A substance for processing that contains at least one platinum    group element, a copper source material containing copper oxide, a    solid carbonaceous reducing agent and flux are all made ready in the    form of powdery or particulate materials, the powdery or particulate    materials are premixed, and the premixed powdery or particulate    material is charged into the electric furnace.-   2. The charged material is heated and melted, whereafter a standing    step is conducted. The standing step is constituted of making the    heat-melted material stand at a temperature of 1200-1500° C. for at    least 5 hours. After the standing step, the molten metal enriched in    platinum group elements is discharged outside the furnace.-   3. The blending ratio of the raw materials is regulated so that the    slag oxides formed in the furnace fall in the content ranges of:    Al₂O₃: 20-40 wt %, SiO₂: 25-40 wt %, CaO: 20-35 wt %, and FeO: 0-35    wt %,-   4. The composition of the oxide raw material is controlled by the    steps of: analyzing and ascertaining beforehand the amount of at    least one among the oxides, preferably all of the oxides, of Al, Si    and Fe contained in the oxide raw material entraining platinum group    elements, and regulating the composition of the flux components    before charging them into the furnace in accordance with the    contents of aforesaid oxides.-   5. The composition of the slag separated from the molten metal is    regulated to:

Al: 10-22 wt %

Si: 10-16 wt %

Ca: 14-22 wt %

Fe: 27 wt % or less (including 0%)

Pt: 10 ppm or less, and

the balance substantially of oxygen.

-   6. As the electric furnace for this method there is used an electric    furnace wherein the furnace wall is constituted of a iron shell, a    continuous sheet of flowing water is formed on the outer surface of    the iron shell, and molten slag forms a solidified self-coating    layer on the inner surface of the iron shell.

These features will now be explained in further detail.

The substance for processing containing platinum group elements(PGM-containing substance) is mixed with flux components (e.g., silica,calcium oxide, calcium carbonate or the like), a carbonaceous reducingagent (e.g., coke powder) and a copper source material (copper or copperoxide) at appropriate ratios and the mixture is charged into theelectric furnace. The ratio of the flux components, while differing withthe composition of the PGM-containing substance, is preferablydetermined so that the vitreous oxides after heating and melting (theelectric furnace slag) has a composition of: Al₂O₃: 20-40 wt %, SiO₂:25-40 wt % and CaO: 20-35 wt %. The slag oxide generated in the furnaceis in the end determined by the PGM-containing substance and fluxcomponents charged into the furnace. The carbonaceous reducing agentdoes not remain in the slag as an oxide and substantially all of thecopper oxide charged as the copper source material is reduced tometallic copper.

The ranges of the constituents of the slag oxide generated in thefurnace can therefore basically be determined by regulating the blendingratio of the PGM-containing substance and the flux components. However,a precondition for this is that the meltdown, reduction reaction,slag-hmetal phase separation and the like proceed in good order. It wasdiscovered that this precondition can be satisfied by, as set out infeature 1 above, preparing the all of the materials to be charged intothe furnace in the form of powdery or particulate materials, premixingthe powdery or particulate materials, and charging the mixture into theelectric furnace. More specifically, the particle diameter of themetallic copper or copper oxide charged into the furnace is preferably0.1 mm or greater to less than 10 mm, and that of 50 wt % of thePGM-containing substance is preferably less than 10 mm. These twocomponents are uniformly mixed with powdery flux and powderycarbonaceous reducing agent and the resulting mixture is charged intothe furnace.

The reducing agent is used chiefly for the purpose of reducing thecopper oxide to metallic copper. The reducing agent used is typicallycoke. However, it is instead possible to use a base metal containing anoble metal or PGM, in which the noble metal or PGM in the base metalcan be simultaneously recovered. Resin, activated carbon and the likecan also be used as the reducing agent. The copper source material isused as a medium for dissolving and incorporating PGM. It can bemetallic copper per se, but copper oxide is also usable.

When the electric furnace is operated using these charged raw materials,the charged raw materials are first heated and melted (meltdown). Themeltdown temperature is 1200° C.-1700° C., preferably 1300° C.-1550° C.At below 1200° C., the slag does not completely melt, so that it becomeshigh in viscosity and the PGM recovery rate decreases. However, atemperature exceeding 1700° C. is undesirable from the viewpoint ofenergy consumption, of course, but also because it may cause damage tothe body of the furnace. Owing to the meltdown, the PGM carrier material(alumina and other oxides) that accounts for almost all of thePGM-containing substance becomes vitreous molten slag that floats, whilethe copper oxide is reduced by the coke or the like to become metalliccopper that, owing to the difference in specific gravity, sinks throughthe slag to form a molten metallic copper layer (molten metal).

After charging of all materials into the furnace has been completed, asealed internal atmosphere is established in the furnace and the chargedmaterial is heated and melted by passage of electric current. As set outin feature 2 above, after the heating and melting, a standing step ispreferably provided for maintaining a temperature of 1200-1500° C. forat least 5 hours. The step of standing for at least 5 hours or longer isestablished between meltdown and molten metal discharge. When the rawmaterial charged into the furnace is prepared as a mixture of powderyand particulate materials and the standing step is established, themolten metal in the electric furnace, which consists almost entirely ofmetallic copper, can capture PGM at a high recovery rate.

When the temperature of the standing step is below 1200° C., the PGMrecovery rate is insufficient regardless of how long the standing periodis prolonged. On the other hand, increasing the temperature above 1500°C. causes damage to the furnace without producing any improvement in thePGM recovery rate. Almost all PGM can be recovered over a standingperiod of at least 5 hours at a temperature in the range of 1200-1500°C. When the standing period is prolonged beyond a certain point, thetendency for the recovery rate to improve reaches saturation. This makesit economically advantageous to complete the standing step after astanding period of, say, around 5 to 10 hours.

PGM is absorbed by the metallic copper as the metallic copper sinksthrough the slag. The recovery rate of the PGM into the metallic copperat this time varies with the material temperature after meltdown and thestanding period. It also varies with the particle diameter of themetallic copper or copper oxide loaded in the furnace, the particlediameter of the PGM-containing substance loaded in the furnace, and thelike. As pointed out above, it is therefore essential to implementappropriate control.

Thus, in the present invention, the standing step is appropriatelycontrolled and the raw materials to be charged into furnace are chargedafter being made particulate and mixed, whereby the molten metal in theelectric furnace, which consists almost entirely of metallic copper, cancapture PGM at a high recovery rate. Although the reason for this is notaltogether clear, it is reasonable to conclude as follows.

At the time point when the carrier material (alumina and other oxides)that accounts for almost all of the PGM-containing substance comes to bemelted together with the flux components it disperses as a slag ofsuitable viscosity. Then, the charged metallic copper or the metalliccopper reduced from copper oxide by the reducing agent disperses intothe slag. This dispersion is particularly good when the metallic copperor copper oxide is incorporated into the mixture as particulate matter.The metallic copper dispersed in the slag absorbs the PGM dispersed andfloating in the slag as its sinks toward the metallic copper of thelower layer under its own weight. Although this process starts frommeltdown, in the case where the temperature is too low (e.g., under1200° C.) during the ensuing standing, the viscosity of the slag risesto the point that the momentum of both the PGM and the metallic copperpresent therein becomes low, so that both stay in the floating state. Onthe other hand, when the temperature during standing is too high (e.g.,above 1500° C.), more than the required amount of heating energy isconsumed, which is uneconomical.

Viewed in this manner, it can be seen that in the standing step it isimportant for the molten metallic copper to be present in slag of anappropriate viscosity in which PGM is dispersed throughout, to be in anappropriate dispersed state in the slag, and to sink slowly through theslag with an appropriate momentum. For achieving a good dispersed state,the charged raw materials must be mixed after being made particulate,while for achieving appropriate viscosity it is important to regulatethe amount added and the composition of the flux components, and also tocontrol the temperature. The standing needs to be continued for a periodof time sufficient for substantially all of the molten metallic copperto sink through the slag. Once reduction of copper oxide proceeds nofurther and the sinking of the molten metallic copper is completed, nomore PGM is taken into the lower layer of molten metal. By achievingthis condition, the method of the present invention is able to absorbPGM into the molten metal at a high recovery rate.

On completion of the standing step, most of the slag is discharged butsome is left in the furnace. Operating time can be shortened if desiredby setting up two electric furnaces side by side and staggering theiroperation so that when one is in the standing state, charging andmelting of the materials is being carried out in the other. It is alsopossible to transfer the heated and melted material in the furnace to aseparate standing furnace and carry out the standing step in it.

As set out in feature 3 above, the blending ratio of the raw materialsis preferably regulated so that the slag oxides formed in the furnacefall in the content ranges of: Al₂O₃: 20-40 wt %, SiO₂: 25-40 wt %, CaO:20-35 wt %, and FeO: 0-35 wt %. For this, as set out in feature 4 above,it is possible to control the composition of the oxide raw materials bythe steps of: analyzing and ascertaining beforehand the amount of atleast one among the oxides of Al, Si and Fe contained in thePGM-containing substance, and regulating the composition of the fluxcomponents before charging them into the furnace in accordance with theascertained contents of said oxides ascertained by the analysis.

More specifically, the PGM-containing substance is crushed intoparticulate matter of a particle size of 5 mm or less before beingcharged into the furnace and a sample for analysis is taken from thiscrushed and mixed raw material to be processed. By regulating thecomposition of flux components (at least one member selected from thegroup consisting of Al₂O₃, SiO₂, CaO and FeO) based on the result of theanalysis, the slag oxides formed in the furnace are regulated to fall inthe content ranges of: Al₂O₃: 20-40 wt %, SiO₂: 25-40 wt %, CaO: 20-35wt %, and FeO: 0-35 wt %.

As a result, the composition of the slag separated from the molten metalis, as set out in feature 5 above, regulated to a composition of Al:10-22 wt %, Si: 10-16 wt %, Ca: 14-22 wt %, Fe: 27 wt % or less(including 0%), Pt: 10 ppm or less, and the balance substantially ofoxygen.

When the slag oxides fall in the aforesaid content ranges of: Al₂O₃:20-40 wt %, SiO₂: 25-40 wt %, CaO: 20-35 wt %, and FeO: 0-35 wt %, thereis obtained a slag that has an appropriate viscosity and readilydisperses and flows, so that in the course of separation owing todifference in specific gravity, the platinum group elements that werepresent in the raw material to be processed are easily absorbed into themolten metallic copper. As a result, the final slag at the end of theprocessing comes to have the foregoing composition of: Al: 10-22 wt %,Si: 10-16 wt %, Ca: 14-22 wt %, Fe: 27 wt % or less (including 0%), Pt:10 ppm or less, and the balance substantially of oxygen.

When the slag generated in the electric furnace deviates from theaforesaid controlled ranges, e.g., when the Al₂O₃ content exceeds 40 wt%, the viscosity of the slag shoots up, and this is believed to be whythe rate of contact between the platinum group elements and moltenmetallic copper reduced from the copper oxide slows to make it easy formolten metallic copper incorporating the platinum group elements to stayafloat in the slag, thus lowering the rate of absorption of the platinumgroup elements into the molten metal. The preferred range of Al₂O₃content is 20-30 wt %.

As set out in feature 6 above, the processes of features 1-5 can all beconducted using an electric furnace having a wall structure formed withthe sheet of flowing water and the self-coating layer explained withreference to FIGS. 1-3. Use of this electric furnace enables recovery ofplatinum group elements at high yield with good operability.

Following the standing step in the electric furnace operation, most ofthe upper layer of slag is discharged and discarded but some is left inthe furnace. Next the molten metal into which platinum group elementswere absorbed is tapped from the electric furnace and transferred to afurnace for oxidizing molten metal (oxidizing furnace) while still inthe molten state, whereafter the platinum group elements are enriched.For this, the melt is oxidized using oxygen gas or an oxygen-containinggas in an oxidizing furnace separate from the electric furnace, therebyseparating it by difference in specific gravity into an oxide layercomposed primarily of copper oxide and molten metal composed primarilyof metallic copper enriched in platinum group elements. The oxidation isconducted by introducing oxygen gas or an oxygen-containing gas whilemaintaining a material temperature of 1100° C.-1600° C., preferably1200° C.-1500° C. At a temperature below 1100° C., the oxidation rate islow, and at a temperature above 1600° C., the furnace body is damaged.Following the oxidation, the upper oxide layer is drained off andseparated by tilting the furnace. Next, the lower molten metal enrichedin platinum group elements is poured out of the furnace and sent to thenext recovery process.

Following completion of the oxidation, the ordinary practice afterdraining off the upper oxide layer is to make up for the amount of thedecrease with fresh molten metal containing absorbed platinum groupelements from the electric furnace, thereby forming a combined melt withthe molten metal remaining in the furnace, and then repeat theoxidation. Preferably, the molten metal is not drained from theoxidizing furnace and sent to the next process, i.e., the platinum groupelement harvesting process, until the platinum group element content ofthe lower molten metal has reached 10% to 75% owing to the repeatedoxidation.

Since the oxide layer drained from the oxidizing furnace is composedalmost entirely of copper oxide, it can be reused as the copper sourcematerial supplied to the electric furnace after being drained off andsolidified by cooling. This also enables recovery of platinum groupelements entrained by the oxide layer. The oxides can bewater-granulated by rapid water cooling from the molten state, namely,can be obtained as particulate matter of a particle diameter of 0.1 mmto 10 mm, which is ideal as a raw material for charging into theelectric furnace.

EXAMPLES Example 1

Spent vehicle exhaust gas purification catalyst containing on averageabout 1200 ppm of Pt, about 300 ppm of Pd and about 90 ppm of Rh(containing on average about 38.5 wt % of Al₂O₃, about 39.6 wt % ofSiO₂, and about 12.5 wt % of MgO) was granulated to a particle diameterof not greater than 10 mm. With 1000 kg of the granulated spent catalystwas mixed 500 kg of CaO and 100 kg of SiO₂ as flux component, 30 kg ofcoke as reducing agent, and 300 kg of copper oxide (about 80% of whichwas particulate matter of a particle diameter of 0.1 mm to 10 mm). Themixture was charged into an electric furnace.

The charged material was melted in the electric furnace by heating toabout 1500° C. After meltdown, it was left to stand under continuedapplication of electric current to hold it at a material temperature ofabout 1400° C. Once every hour, part of the upper layer of slag wasdrained from the side of electric furnace and solidified by cooling.This operation was continued for 20 hours after meltdown. The PGM in theslag sampled hourly was analyzed. The results are shown in Table 1.

From the results in Table 1 it can be seen that at this holdingtemperature considerable amounts of Pt, Pd and Rh remained in the slagwhen the standing period was within 5 hours, but that the amounts becamevery low when the standing period exceeded 5 hours, with the tendency todecrease substantially stopping at around 8 hours.

Example 2

Example 1 was repeated except that the holding temperature was set at1200° C. and the standing period at 5 hours. The PGM in the slag wasanalyzed as in Example 1. As shown in Table 1, the results were Pt: 0.9ppm, Pd: 0.2 ppm and Rh: 0.1 ppm.

Example 3

Example 1 was repeated except that the holding temperature was set at1300° C. and the standing period at 5 hours. The PGM in the slag wasanalyzed as in Example 1. As shown in Table 1, the results were Pt: 0.7ppm, Pd: 0.1 ppm and Rh: 0.1 ppm.

Comparative Example 1

Example 1 was repeated except that the holding temperature was set at1100° C. and the standing period at 5 hours. The PGM in the slag wasanalyzed as in Example 1. As shown in Table 1, the results were Pt: 2.5ppm, Pd: 0.9 ppm and Rh: 0.2 ppm. At a holding temperature under 1200°C., the PGM could not migrate sufficiently from the slag to the metal.

Comparative Example 2

Example 1 was repeated except that the holding temperature was set at1550° C. and the standing period at 5 hours. The PGM in the slag wasanalyzed as in Example 1. As shown in Table 1, the results were Pt: 1.5ppm, Pd: 0.4 ppm and Rh: 0.1 ppm. At a holding temperature above 1500°C., the PGM could not migrate sufficiently from the slag to the metal.TABLE 1 Standing Standing Amount of PGM period temp. Loss into slag (hr)(° C.) Pt Pd Rh Remark 0 1400 9.7 2.5 1.0 Example 1 1 1400 5.2 1.5. 0.32 1400 3.8 1.2 0.3 3 1400 2.5 1.0 0.2 4 1400 1.9 0.6 0.2 5 1400 0.9 0.2<0.1 6 1400 0.8 0.2 <0.1 8 1400 0.8 0.1 <0.1 10 1400 0.7 0.1 <0.1 201400 0.6 <0.1 <0.1 5 1100 2.5 0.9 0.2 Comparative Example 1 5 1200 0.90.2 <0.1 Example 2 5 1300 0.7 0.1 <0.1 Example 3 5 1550 1.5 0.4 0.1Comparative Example 2

Example 4

The molten metal in the electric furnace after conducting an operationof 8 hours of standing at 1400° C. in accordance with Example 1 wastapped from the electric furnace and led into a heated oxidizingfurnace. Oxygen-enriched air of an oxygen concentration of 40% was blownonto the surface of the molten metal in the oxidizing furnace until anapproximately 1-cm thick oxidized layer had formed on the surface of themelt. At this point, the furnace was tilted to discharge the layer ofoxide from the furnace into a water tank through which a large quantityof water was being passed. The furnace was returned to its originalorientation and the operation of blowing oxygen-enriched air onto themelt surface to form an oxidized layer and then discharging the layerinto the water tank when it had reached a thickness of approximately 1cm was repeated. This rapid cooling in the water formed water-crushedparticles of a particle diameter of 10 mm or less. The result could beused as part of the copper oxide component of the raw material chargedinto the electric furnace.

To the molten metal remaining in the oxidizing furnace after dischargeof the oxide layer was added with molten metal obtained by processingcorresponding to that on the electric furnace side in Example 2 andoxygen-enriched air was similarly blown onto the resulting surface. Theoperation of discharging the oxide layer from the oxidizing furnace whenit had reached a thickness of approximately 1 cm was repeated twice. Allof the melt remaining after this operation was then discharged from theoxidizing furnace and solidified by water cooling. About 10.5 kg ofmetallic copper was obtained. The PGM content of this metallic copperwas: Pt: about 22 wt %, Pd: about 5.5 wt %, Rh: about 1.5 wt %.

Example 5

Spent honeycomb-shaped automotive exhaust gas purification catalyst,1000 kg, was loaded on a conveyor and supplied to a jaw crusher at thefirst stage and a double-roll crusher at the second stage, therebycrushing the whole amount to a particle diameter of 5 mm or less. Thewhole amount of the crushed raw material to be processed was passedthrough two three-stage rifflers (at a reduction of ½×½×½=⅛ in eachriffler) and 15.5 kg of a 1/64 riffled representative sample was taken.The remaining bulk was stored in a silo. The whole amount of therepresentative sample was dried and measured for water content (=0.5 wt%). Next, the whole amount was pulverized to under 100 mesh in apulverizer and mixed in a V-blender, whereafter an analysis sample, 100g, was colledted using a rotary 1/12 riffler.

The oxides of the analysis sample were analyzed using a fluorescentX-ray analyzer and were found to consist of Al₂O₃: 40.5 wt %, SiO₂: 41.6wt %, MgO: 11.5 wt %, and FeO: 1.5 wt %.

The desired makeup of the slag generated in the electric furnace wasdefined as: Al₂O₃: 22.3 wt %, SiO₂: 28.5 wt %, CaO: 28.1 wt % and FeO:12.1 wt %. With the aim of using the aforesaid analysis values as thebasis for achieving this desired makeup, there was weighed out 984.5 kgof the bulk stored in the silo and, as flux components, 500 kg of CaO,100 kg of SiO₂ and 200 kg of FeO. In addition, 30 kg of coke as reducingagent and 300 kg of copper oxide as copper source material (about 80% ofwhich was particulate matter of a particle diameter of 0.1 mm to 10 mm)were weighed out and the total amounts of the four types of materialwere mixed.

The mixture was charged into an electric furnace and melted at atemperature of about 1350° C. After meltdown, the melt was left to standfor about 5 hours at a temperature of 1250 to 1300° C. The upper layerof slag oxides was then discharged from the side of the electric furnaceand solidified by cooling. Analysis of the platinum group elements inthe slag showed it to contain: Pt=0.7 ppm, Pd=0.1 ppm and Rh=0.1 ppm,i.e., loss of platinum group elements into the slag was very slight.

Analysis of the slag oxide components showed it to contain: Al₂O₃: 21.5wt %, SiO₂: 29.2 wt %, CaO: 27.9 wt % and FeO: 11.8 wt %. All of thesevalues are within +1.0 wt % of the aforesaid desired makeup values.

Reference Example 1

From 1000 kg of large and small broken chunks of spent honeycomb-shapedautomotive exhaust gas purification catalyst (converter pieces)contained in two flexible containers, 15 kg was randomly selected as arepresentative sample. The whole amount of the representative sample wasdried, measured for water content (=0.8 wt %), and crushed in a jawcrusher. The whole amount of the crushed material was pulverized tounder 100 mesh in a pulverizer and mixed in a V-blender, whereafter ananalysis sample, 100 g, was colledted using rotary 1/12 riffler.

The oxides of the analysis sample were analyzed using a fluorescentX-ray analyzer and were found to consist of Al₂O₃: 37.8 wt %, SiO₂: 43.1wt %, MgO: 12.3 wt %, and FeO: 1.2 wt %.

The desired makeup of the slag generated in the electric furnace wasdefined as: Al₂O₃: 22.0 wt %, SiO₂: 25.1 wt %, CaO: 29.1 wt % and FeO:12.6 wt %. With the aim of using the aforesaid analysis values as thebasis for achieving this desired makeup, there was weighed out 985 kg ofthe chunks of spent honeycomb-shaped automotive exhaust gas purificationcatalyst (converter pieces) and, as flux components, 500 kg of CaO and200 kg of FeO. In addition, 30 kg of coke as reducing agent and 300 kgof copper oxide as copper source material (about 80% of which wasparticulate matter of a particle diameter of 0.1 mm to 10 mm) wereweight out. The materials were charged into an electric furnace andmelted at 1350° C.

The charged material was melted in the electric furnace at about 1350°C. After meltdown, the melt was left to stand for about 5 hours at atemperature of 1250 to 1300° C. The upper layer of slag oxides was thendischarged from the side of the electric furnace and solidified bycooling. Analysis of the platinum group elements in the slag showed itto contain: Pt=1.8 ppm, Pd=0.4 ppm and Rh=0.2 ppm, i.e., loss ofplatinum group elements into the slag was greater than in Example 5.

Analysis of the slag oxide components showed it to contain: Al₂O₃: 23.5wt %, SiO₂: 22.1 wt %, CaO: 29.5 wt % and FeO: 11.8 wt %. The contentvalues for Al₂O₃ and SiO₂ are more than 1.5% higher than the aforesaiddesired makeup values.

Example 6

The molten metal containing platinum group elements in the electricfurnace after conducting an operation according to Example 1 was tappedfrom the lower part of the electric furnace and led to a heatedoxidizing furnace. Oxygen-enriched air of an oxygen concentration of 40%was blown onto the surface of the molten metal in the oxidizing furnaceuntil an approximately 1-cm thick oxidized layer had formed on thesurface of the melt. At this point, the furnace was tilted to dischargethe layer of oxide from the furnace into a water tank through which alarge quantity of water was being passed.

The operation of blowing oxygen-enriched air onto the surface of themolten metal in the oxidizing furnace to form an oxidized layer and thendischarging the layer into the water tank when it had reached athickness of approximately 1 cm was repeated. After five repetitions ofthis operation, the molten metal containing platinum group elements inthe electric furnace after conducting an operation according toComparative Example 1 was drawn from the lower part of the electricfurnace and led to a heated oxidizing furnace, thereby forming acombined melt. The operation of blowing oxygen-enriched air onto themolten metal in the furnace, discharging the generated oxides from thefurnace and water-cooling them was repeated. After this processing allof the obtained molten metal was discharged from the oxidizing furnaceand solidified by water cooling. Analysis of the result showed that theamount of metallic copper was 5.4 kg, and that the platinum groupelement contents were: Pt: 21.3 wt %, Pd: 6.7 wt % and Rh: 1.4 wt %.

1. A method for recovering platinum group elements comprising: chargingan oxide raw material entraining platinum group elements, copper oxide,solid carbonaceous reducing agent and flux into a closed electricfurnace equipped with electrodes for passing electric current throughand heating material charged in the furnace and a roof for substantiallyshielding material charged in the furnace from the exterior atmosphere,melting and reducing the charged material by the electric currentheating with the electrode, thereby forming a layer of molten metalunder a molten slag layer and enriching platinum group elements in thismolten metal, which method is characterized in that: the oxide rawmaterial, copper oxide, solid carbonaceous reducing agent and flux areall prepared as powdery or particulate materials; these powdery orparticulate materials are premixed and then charged into the electricfurnace; and the molten metal enriched in platinum group elements isdischarged outside the furnace after undergoing a standing step in whichthe heat-melted material in the furnace is allowed to stand at atemperature of 1200-1500° C. for at least 5 hours.
 2. A method forrecovering platinum group elements according to claim 1, wherein theslag oxides formed in the furnace are controlled to the content rangesof: Al₂O₃: 20-40 wt %, SiO₂: 25-40 wt %, CaO: 20-35 wt %, and FeO: 0-35wt %.
 3. A method for recovering platinum group elements according toclaim 2, wherein the content ranges of the slag oxides are controlled bythe steps of: analyzing and ascertaining beforehand the amount of atleast one among the oxides of Al, Si and Fe contained in the oxide rawmaterial entraining platinum group elements, and regulating thecomposition of the flux components before charging them into the furnacein accordance with the contents of said oxides.
 4. A method forrecovering platinum group elements according to claim 3, wherein theflux contains at least one component selected from the group consistingof Al₂O₃, SiO₂, CaO and FeO.
 5. A method for recovering platinum groupelements according to claim 1, wherein the composition of the moltenslag is regulated to: Al: 10-22 wt % Si: 10-16 wt % Ca: 14-22 wt % Fe:27 wt % or less (including 0%) Pt: 10 ppm or less, and the balancesubstantially of oxygen.
 6. A method for recovering platinum groupelements according to claim 1, wherein the molten metal discharged fromthe furnace is transferred to and oxidized in another furnace andseparated by difference in specific gravity into an oxide layer composedmainly of copper oxide and molten metal composed mainly of metalliccopper in which platinum group elements are enriched.
 7. A method forrecovering platinum group elements according to claim 1, wherein themolten metal discharged from the furnace is transferred to and oxidizedin another furnace and separated by difference in specific gravity intoan oxide layer composed mainly of copper oxide and molten metal composedmainly of metallic copper in which platinum group elements are enriched,the oxide layer composed mainly of copper oxide obtained from thedischarged molten metal being reused as the copper oxide for thecharging step.
 8. A method for recovering platinum group elementscomprising: charging an oxide raw material entraining platinum groupelements, copper oxide, solid carbonaceous reducing agent and flux intoa closed electric furnace equipped with electrodes for passing electriccurrent through and heating material charged into the furnace and a rooffor substantially shielding material charged into the furnace from theexterior atmosphere, melting and reducing the charged material by theelectric current heating with the electrode, thereby forming a layer ofmolten metal under a molten slag layer and enriching platinum groupelements in this molten metal, which method is characterized in that: asthe electric furnace there is used an electric furnace whose wallenclosing up to a height level of the molten slag layer generated in thefurnace is constituted of an iron shell and a sheet of flowing water isformed on the outside of the iron shell to descend in contact with itsouter surface.
 9. A method for recovering platinum group elementsaccording to claim 8, wherein a self-coating layer consisting of asolidified layer of fused substances in the furnace is produced on theinner surface of the iron shell.
 10. A method for recovering platinumgroup elements according to claim 8, wherein the sheet of flowing waterdescending in contact with the outer surface of the iron shell is formedby evenly distributing water of a prescribed water head pressure from aheader installed outward of an upper part of the furnace wall toward thewhole outer periphery of the iron shell at the same height level.
 11. Amethod for recovering platinum group elements according to claim 9,wherein the sheet of flowing water descending in contact with the outersurface of the iron shell is formed by evenly distributing water of aprescribed water head pressure from a header installed outward of anupper part of the furnace wall toward the whole outer periphery of theiron shell at the same height level.